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By Feng Cai

"As a new method to support gate roadways, rock bolting has been widely used in the big coal enterprises for its superiority. The disturbed stress field of surrounding rock and anchor pre-stress field were divided into 3 areas in consideration of s1, s3. The surrounding rock failure form and characteristic were analyzed in combination with parameter f, the failure mechanism was explained using the slip line theory, furthermore, the rock bolting mechanism and its arrangement were studied. After roadway excavation, the tensile stress area appeared at the surface of the surrounding rock. The failure type changes from shallow tension-shear form into deep compression-shear form and the failure shape is an “incense burner”, surrounding by high risk failure area in butterfly shape. The maximum principle stress in roof appears discontinuous distribution because of layers separation. The failure type of surrounding rock is intensity-stress environment-weak plane control model. The internal reason of roadway macroscopic failure is the displacement between the surface plastic area and the deep elastic region along the tangential direction of slip line. In condition of rock bolt support, there is one-way tensile stress area in small range at the surface of surrounding rock, inside is the bidirectional compressive zone in uniform distribution form, stress environment is improved by the superposition of anchor pre-stress and disturbed stress field, anchor rock bolt arrangement and pre-tightening force distribution is decided according to the distribution of f parameter. Finally, supporting method and parameter reasonability was proved by the roadway deformation monitoring, damage depth drilling peep test in the 8101 transportation roadway in a mine.Introduction The stability of roadway support is very important to the safety and sustainable development of coal mine. In recent years, the focus of China's coal mining industry began to move to West China, the thick-extra thick coal seam is developed in the West. The latest technology and equipment have been put in to service, the occurrence condition of thick coal seam and the increasing of productivity lead to the cross section of the roadway increasing continually, hence, it is more difficult to control the surrounding rock. At present, the most common way to support the roadway is rock bolt (cable) support (Wang 2007); Scholars have been deeply studied on the mechanism of rock bolt. It is reported (Wang et al 2012, Gao 2007, Zhang and Yuan 2006) that the influence of rock bolt on physical and mechanical parameters of coal and rock mass in the anchorage zone, It is believed that the rock bolt can improve the peak strength of surrounding rock, especially the post peak strength, and the failure mode of rock mass under the rock bolt support is analyzed. It is reported (Wang et al 2014, Xiao et al 2013, Li et al 2007, Yan et al 2012), according to different geological conditions, the failure mechanism of surrounding rock and supporting technology of the whole coal roadway were studied. Focus on the failure mechanism of roadway surrounding rock, different forms of high strength support methods were proposed, which achieved good results. It is reported (Wang et al 2012a, Zhang et al 2011, Wang et al 2012b, Kang et al 2010, Zhang et al 2010, Kang et al 2007) that the supporting mechanism of pre-stressed anchor was analyzed from the point of the distribution of the maximum principal stress in the pre-stressed anchor field, and the influence of the rock bolt support parameters and surrounding rock characteristics on the supporting effects were studied. Based on the latest research results, this paper analyzed the failure mechanism of surrounding rock of coal roadway, proposed the high strength anchor net cable combined supporting technology based on the feature of stress zoning distribution of surrounding rock, to provide reference for the coal roadway rock bolt (cable) support technology."

Jan 1, 2017

By J. M. Applegate, C. L. Morgan

Introduction In the same way that a blueprint communicates informa¬tion about a structure, a systems architecture should provide information that guides the construction and implementation of a computing system. Taking the blueprint analogy a step further, a plot plan can also show where, in the "bigger picture," a particular design fits (if in fact it does) and what a builder needs to know about the environment in which he or she will operate. In the context of business systems, a meaningful architecture should convey the following: • the primary purpose of computer-based systems rela¬tive to particular tasks or operations, • the relationship of the systems to each other and an indication of their interaction and dependencies, • the relationship of the systems relative to the organiza¬tion and its objectives, • indications of where IS/IT resources can best be allo¬cated to support the organization and • the overall context in which new systems and technolo¬gies can be assessed for both their impacts on current systems and on the organization itself. In theory, a well-designed systems architecture should allow an organization to focus its resources on those aspects of its business that matter most, thus, leveraging investments in information technology to a competitive advantage. In addition, the way that a computing systems architecture is represented should be un¬derstandable to non-IT staff, and it should be eas¬ily recognized by the people it supports. This pa¬per attempts to show that a meaningful systems model exists and that its applica¬tion to mining can achieve the desired results. The following discus¬sion is based on the au¬thors interpretation of a concept referred to as the computer integrated manu¬acturing (CIM) model, as developed by Allen-Bradley. An inherent strength of the CIM approach is that the model combines aspects of business-reporting structures with the manufacturing process itself. What results is a systems frame¬work that is aligned with relevant business processes and is, therefore, easier for non-IS/IT personnel to interpret. The successful application of the CIM model can be seen in a large number of both discrete and process-manufacturing facilities throughout the United States. This paper provides a brief overview of the CIM model and then describes how the model can be applied at a conceptual level to the mining industry. Finally, this paper proposes an extension to the CIM model that the authors believe will make it easier to understand and apply. CIM model As shown on Fig. 1, the CIM model begins with the most detailed levels of process equipment and works its way up to the highest level of consolidated financial reporting. Systems at each level are characterized by common attributes, users, time frames, etc., providing a clear picture of their interac¬tion. Further analysis of the relationships between systems and between levels provides insight into the opportunities for integration. Regardless of the industry or processes involved, the CIM model provides a framework in which the characteristics of systems and the data they manage can be predicted based on the nature of the business function they support. In general, as data moves upward from the lower rungs of the CIM model, it changes from de¬scribing the physical na¬ture of a process to describ¬ing its financial aspects (e.g., from productivity to cost). The time frame to which data applies also transitions from a very short duration, perhaps only milliseconds, to financial or regulatory reporting time frames that may span years. In a simi¬

Jan 1, 2000

By Herschel H. Potter

While preparing this paper, I soon realized I had a problem in defining what injuries should be classified as "personal." Admittedly, in some cases, falls of persons, sprains and strains, pulling and pushing injuries, and back injuries may have been caused by the individual. But, I believe the number of injuries directly attributed to an individual's carelessness is relatively small in comparison to the total number of injuries in those categories. For discussion purposes, the injuries most often considered to be personal, that is, caused by individuals are those classified as back injuries. My paper will be devoted to back injuries because this category represents the greatest number of injuries occurring in coal mining. A man's best friends are his wife, his back, and his dog. The back, however, has a reputation of not being as faithful as his other two friends.1 "Industrial back injury is one of the major occupational health problems plaguing industry to- day. According to national surveys, back injury accounts for more lost time than any other occupational injury and costs companies thousands of dollars in compensation, medical fees and over- time pay."2 A study conducted by USA Today found that back ailments afflict an estimated 75 million Americans, have totally disabled about 2.5 million adults, lead to as many as 200,000 required surgeries each year, are a leading cause of absenteeism and disability, are the second leading cause of hospitalization, are responsible for 18 million doctor visits a year and cost about 93 million work-days annually. The cost to industry is estimated at $14 billion a year.3 During the past 30 years, a great deal of accident prevention education has been directed toward the reduction of back injuries. However, this effort has not produced any great change in the incidence of the injury. Back injuries should be of the utmost concern to everyone here today. I don't need to tell you what the incidence rate is for back injuries, nor will I use any statistics to inform you of the incidence rate for the coal mine industry or any specific mine. It is only important to know that the number of back injuries is excessive. Back in- juries resulting from lifting have been and continue to be one of the most constant and undiminishing problems in the occupational environment. The fact is, we use our backs to per- form work that the back was not designed to do. Historically, manual materials handling related accidents have consistently accounted for the largest percentage of all industrial accidents. In fact, over 25 percent of all industrial accidents are related to handling of materials.4 In underground coal mining operations, material handling accidents account for even larger percentage of lost time injuries. In 1983, for example, they accounted for 34 percent of all lost time injuries and 32.6 percent of all non-lost time injuries. 5 I want to use one example of a task that will illustrate what I have just said: the building of a concrete block stopping. Each solid concrete block will weigh from 60 to 70 pounds. To build a stop- ping 6 feet high and 20 feet wide will take approximately 80 blocks. On the average, these 80 blocks will be handled at least four times from loading into the conveyance for transportation to being placed while constructing the stopping. This generally requires repetitive lifting from ground level. The next three slides demonstrate the lifts

Jan 1, 1986

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By Patrick J. Maessen, Bart H. M. Hendrix

"In the Netherlands, the Dutch Tunnel Standard has been developed for the execution of road tunnel projects. The Dutch Tunnel Standard is based on the Dutch Tunnel Law (WARVW) which sets the minimum safety standard for road tunnels in the Netherlands. Uniformity of tunnels and the uniform perception of a tunnel by its users leads to a higher safety level. Therefore, several items in the Dutch Tunnel Standard are specifically developed to ensure safety in emergency situations through the plane of human factors. Taking human factors into account is important in regular traffic situations, but in an emergency situation, this can be vital. In this paper, the development and contents of the Dutch Tunnel Standard are explained in the first part and in the second part, the systems related to human factors are described. DUTCH TUNNEL LAW AND THE NEED FOR THE DUTCH TUNNEL STANDARD In the 90’s of last century and the early 2000’s, several serious accidents occurred in tunnels in Europe. These accidents involved fire and led to many casualties. After these accidents, the European Union decided to change the regulations for tunnel safety, leading to the EU directive 2004/54/EG. The primary goal of this directive is, that casualties are minimized in case of a calamity in a tunnel (e.g. fire). In the Netherlands, the EU directive led to the development of the tunnel law and a tunnel regulation (in 2006) to prescribe the safety features required for a safe tunnel. The new Dutch tunnel law – cited in the Building Act – was meant to make tunnel design easier – but resulted at first in the opposite effect. Tunnel design in the Netherlands and design standards of Rijkswaterstaat Rijkswaterstaat – the executing organisation of the Dutch Ministry of Infrastructure and the Environment – is the owner of the tunnels in the Dutch national motorway network, thus owner of the majority of the road tunnels in the Netherlands. In the early years of tunnel building in the Dutch national highway network (approximately 1950 – 1970), Rijkswaterstaat was not only owner, but also designer of its own assets. Most Dutch tunnels are submersed tunnels, because these type of tunnels provide a cost-effective way to make river crossings. Another type of tunnel that is used, is a TBM tunnel. Because of the geological conditions in the Netherlands (merely clay/peat and sand), there are no NATM or drill-and-blast tunnels in the Netherlands. Since the last decades, land tunnels (covered roads) are becoming more common to provide for multiple use of surface space and to meet more modern environmental regulations as traffic increases near densely populated areas. Tunnel lengths in the Netherlands vary from 500 m to 1.5 km for river crossings up to 6 km for the Western Scheldt (bored) tunnel. The lengths of land tunnels vary from 500 m up to 3 km. With these increasing lengths, some safety topics and design topics arise, topics that were non-issues in the past, when tunnels were only used as river crossings."

Jan 1, 2016

By Charles H. Tiernon

Separation of particles on a table deck* results from differences in specific gravity, plus a number of other interrelated phenomena: size of particles; shape of particles; stratification (consolidation trickling); and hindered settling. The effective use of these factors may be controlled and aided by the following table adjustments: differential action of the drive or head motion; speed of motion; length of stroke; side tilt; end elevation; size of deck area; shape of deck; character of deck surface; riffle design; amount and distribution of wash water; feed water-to-solids ratio; feed rate; and feed uniformity. The typical concentrating table used in the US coal industry is essentially a rhombohedral-shaped deck supported by a suitable subframe. It generally slopes in the direction of the short axis, tilts upward at the refuse end of the long axis, and is differentially reciprocated in the direction of the long axis. This head motion tends to move the particles longitudinally toward the end of the table. At right angles to this movement, wash water (distributed along the high side of the table) tends to move particles down the slope parallel to the short axis of the table. Hence, the natural path of a particle on a concentrating table is from the feed box corner to the diagonally opposite corner of the deck (Fig. 1). There is a natural tendency for heavier particles to sink to the deck of the table and for lighter particles to stay nearer the surface. As a consequence, the heavier particles are carried further along the longitudinal axis since, being on or near the surface of the deck, they are more affected by the table's reciprocating action. The lighter particles, being at or near the top of the moving bed, travel slower in a longitudinal direction and thus separate from the heavy particles. Wash water and the diluting water in the feed cause lighter gravity particles to move down the short axis of the table. The feed material then fans out from the feed box, where it enters onto the table with heavy gravity particles moving in one direction and lighter particles moving in another. By properly adjusting all of the previously mentioned factors, a clear-cut separation is possible between high- and low-gravity particles, or between refuse and coal. Riffles are of great importance to tabling. They create considerably more capacity than is possible with smooth decks, and they are responsible for better separation accuracy. Particles in their downward movement along the deck's slope are delayed behind the riffles. Hindered settling, consolidation trickling, and other phenomena occur as the particles tumble over the riffles and stratification ensues. The heavier particles settle to the deck of the table and are trapped behind the riffles, thus insuring their positive movement toward the end of the table. Since stratification and separation of particles are not complete as a result of any one riffle, a series of riffles are used. The cycle of stratification and hindered settling is repeated from riffle to riffle. The result is purer products as particles fan out and progress forward and downward over the table. The smallest, heaviest particles stratify downward to the surface of the deck more quickly and are carried out by table movement toward the refuse end at a faster rate than coarse, heavy particles. Light gravity, larger pieces ride on the top layer of particles and flow down the deck's slope as a result of the cross flow of wash water at right

Jan 8, 1980

By D. K. Adjei, G. Ofori-Sarpong, R. K. Amankwah

"Mercury (Hg) is toxic and has long-term effect on the muscle tissues, the brain and the central nervous system functions due to its bioaccumulation and biotransformation effect. Hg pollution around the world is high and this has attracted much global concern such as the Minamata Convention on Mercury, held in 2013. For several years, mercury has been the chemical of choice for Artisanal and Small-Scale Mining (ASM) gold extraction, accounting for about one-third of the global mercury consumption. A greater percentage of this ends up in the environment through open burning of amalgam, spillage and indiscriminate disposal of waste. To break this wall of mercury pollution, it is very important for the mercury used to be recycled through the use of mercury retort units. This brings into the limelight, the Lantern Retort which tries to synergise the high conductivity of steel and the see-through effect of glass, thus, improving the amalgamation process. A research conducted on improving the original Lantern Retort arrived at an appropriate distillation length of 5.08 cm which was used in redesigning a more efficient retort. The improved Lantern Retort recovered about 98% of the mercury used. INTRODUCTION Artisanal and Small-scale mining (ASM) is an informal mining activities which employ quite simple tools such as pickaxes, sluice boxes, shovels, chisels, metal bars, mattocks and water pumps for mining and processing of gold ores. The artisanal and small-scale mining (ASM) sector has served as source of employment for many people in developing countries including Ghana (Aryee, 2001; Hilson and Potter, 2003; Amankwah and Anim-Sackey, 2004; Amankwah et al., 2010) and has provided a substantial amount of foreign exchange through its activities (Hilson, 2001; Styles et al., 2010). Hentschel et al. in 2003 estimated that about thirteen (13) million people are said to be engaged directly in the small-scale mining activities worldwide. However, an estimated 100 million people are said to depend on this sector for their livelihood mainly in developing countries. In China alone, a projected number between 3 and 15 million people are engaged in artisanal and small-scale mining activities (Hentschel et al., 2003), whilst in Ghana, it is estimated that more than 1,000,000 people are employed directly or indirectly in the artisanal and small-scale mining activities, with 10% being women (Amankwah and Anim-Sackey, 2004; Amankwah and Ofori-Sarpong, 2014). In Ghana, ASM gold production has grown from a low of 2.2% in 1989 to a high of 34% in 2014. This figure is estimated to increase with proper documentation. This indicates that the sector has made significant socioeconomic impact on many individuals and communities as a whole. Notwithstanding the above positive contributions, the sector has also contributed to negative impacts on human, aquatic life and the environment in general due to indiscriminate mercury emission. This ‘pollution wall’ needs to be broken, and several interventions have been put in place including the use of retort units like the lantern retort (Adjei et al., 2014; Amankwah and Ofori-Sarpong, 2014)."

Jan 1, 2017

By W. L. Wilson

This article could be entitled "The Rediscovery of the Windfall Mine," since it was originally discovered early in this century. The "Rediscovery" was made by personnel of Idaho Mining Corp. during the course of an investigation of the Windfall mine beginning in 1967. There had been historic production from the Windfall mine sporadically from 1908 until 1951 and there was present during much of this time a mill that vat leached crushed ore. Mining was underground to a depth of 91 m (300 ft). Prior to Idaho's activities, the mine had been inactive since the early 1950s. Idaho Mining Corp. was an "exploration company," privately owned and based in Reno, NV. The writer was at that time manager of exploration for Idaho Mining. The company held interests in the Cortez Gold mine, then being readied for production, and had conducted exploration for various mineral commodities in Nevada for several years. Idaho's attention was first directed to the Windfall mine in the Eureka District, Eureka County, Nevada (see Fig. 1, Nevada map) by a search of published geologic literature of the state of Nevada. An attempt was made to acquire a library consisting of all of the published geo-data for the state of Nevada as well as to obtain as much of the unpublished material as could be acquired. A study of these data, and in particular Tom Nolan's Professional Paper 406 (1962), indicated that the Windfall mine was anomalous for the Eureka District in that it contained deposits of disseminated gold. This source and other sources indicated that the Windfall had produced from underground operations approximately 59 kt (65,000 tons) grading 12.6 g/t (0.368 oz per ton) gold ore that was processed in a vat leaching operation and that the gold was "invisible" to the naked eye and could not be panned. It was ascertained that there were no active operations near the mine and a geologic reconnaissance was made and a few samples were taken to attempt to determine if the mineralization was indeed disseminated. The initial group of samples around the mine area averaged 1.25 ppm [11.5 g/t gold (0.036 oz per ton)] and the mineralization did appear to disseminate into the wallrock. The next step involved researching the ownership of property in the immediate area of the mine, and [ ] a property map was constructed that showed six or seven different ownerships of patented and unpatented claims, as well as some open ground, in the target area. Negotiations were commenced with the claim owners, ultimately resulting in the acquisition of all of the desired property, although some of

Jan 1, 1990

By David M. Stearns

Introduction Millions of tons of coal are shipped by rail each year in the US. Loading those trains efficiently is a topic being studied by coal producers and railroads. Alternatives range from volumetric loadouts to systems that weigh the product, using either belt conveyor scales, track scales, or weighbins. Each system has its proper application and each has pros and cons, depending on the volume, nature, and layout of the specific loading facility. In recent years, a batch weighing system has been developed that offers significant benefits to higher volume operations shipping by a unit train. The Unitrain Loadout System (ULS) features a precision electronic batch weighing system married to a mechanical loadout system by a microprocessor-based digital control system. A batch weighing loadout could be designed using analog electronics or even a mechanical weighbin. But it is the adaption and use of a microprocessor that provides the primary benefits of a unit train loading facility. Those benefits are accurate weight determination, optimum car use, overload prevention, speed in loading, optimum labor use, and documentation of the loading process. Unitrain History The first ULS was erected for iron ore applications in Canada during the 1960s. The first domestic system on coal was installed at Pittsburg and Midway Coal Co.'s McKinley mine at Gallup, NM, in 1978. At present, there are more than 20 Unitrain Loadout Systems in operation worldwide. They are loading minerals, concentrates, and coal. Eleven systems are presently in service or under construction in the US. In the last several years, a number of comparable systems have also been installed. As a result, the concept has been well proven and is now commercially available from several sources. Unitrain Operation The basic operation of a ULS involves loading a unit train with individually weighed batches of coal while it moves slowly under a loadout tower. Typical coal systems are designed to load at rates of 2.7 to 6.3 kt/h (3000 to 7000 stph). The contents of each car can be accurately weighed. This meets the requirements of the National Bureau of Standards Handbook No. 44 for static weighing. Also, cars can be loaded within very close tolerances. This optimizes railcar usage and prevents overloads. Principal components of the loadout system are the main feed conveyor, the surge bin, surge bin gates, weighbin, weighbin gate, flood loading chute, control room, calibration weights, and hydraulic and electrical systems (Fig. 1). Main feed conveyor: Coal is delivered from a storage area by the main feed conveyor. The conveyor fills the surge bin. It should be capable of conveying coal at the same rate at which it is being loaded. Usually the feed conveyor is equipped with a conveyor scale that can display rate in the control room and control feed to the belt. The surge bin will typically store 227 to 272 t (250 to 300 st) of coal. If the loadout is installed in an over-the-track silo, there is no need for a surge bin. Four double-bladed, hydraulically actuated gates are incorporated into the bottom of the surge bin. These gates open to fill the weighbin. Automatic pre-act points are selected by the control system. They shut these gates as the desired weight is approached. The final amount of coal is added by a single gate that closes. This results in a weighed batch of coal within ± 0.5% of the selected optimum load per railcar. Thus, if loading 91-t (100-st) capacity cars, the system is set to batch up 90.2 t (99.5 st) batches. Each car is expected to be loaded between 90 and 91 t (99 and 100 st). Weighbin: The weighbin is next in the flow scheme. Typically, it is

Jan 3, 1985

By Norman L. Kotraba

In Electric Arc Furnace (EAF) steelmaking, dust emissions are collected in a baghouse to prevent their escape to the Environment. This fine dust, predominantly iron oxide, also contains oxides of various other metals introduced into the EAF from the scrap charge. For example, mixed or galvanized scrap brings with it zinc, lead, and small quantities of cadmium which preferentially accumulate in the collected off-gas dust. Stainless steel scrap brings chromium into the charge, some of which also makes its way into the baghouse dust as hexavalent chromium-containing oxides. Because many of the contained heavy metal oxides are water-leachable, the U. S. Environmental Protection Agency has listed EAF dust as a hazardous waste which, as of August 8, 1988, cannot be 1andfi11ed without first being treated. Moreover, beginning August 8, 1991, all dusts with zinc concentrations of 15% or greater will be required to undergo high temperature thermal treatment to remove the heavy metals prior to any further disposal. This deadline represents several EPA-granted time extensions to the original requirements; extensions which were granted to allow for further development of process technologies.

Jan 1, 1991

By Todd S. Harris

Industrial minerals or non-metallic minerals are loosely used terms applied to a diverse assortment of materials which, in many cases, have nothing more in common other than the fact that they are neither metals nor fuel minerals. To help bring order to this otherwise random assortment of minerals, we have grouped these materials into two main categories; chemical and physical minerals, and nine sub-categories based on end-use applications as shown in Table 1. Chemical minerals are products which are principally used for the specific chemical elements they contain, whereas physical minerals are products which are used to meet specific performance specifications. The latter part of the 1970?s, in particular 1978, were considered peak years in terms of production and sales for many industries, By using 1978 as a base year, we can evaluate wherethese markets stand today after the recession of the early 1980's and the recovery which began several years ago. From this historical overview and appraisal of the current situation, various assumptions can be drawn as to which markets are expected to perform well into the 1990's.

Jan 1, 1987

By Daniel T. Hegyan

The employment of continuous rotary vacuum disc filters to filter refuse from the fine coal washing circuits has been increasing as it has become more important to close the water circuits and eliminate sludge holding ponds. This paper will discuss our past experience filtering refuse, effects of solids concentration in the filter feed, use of polymer flocculation aids, effect of flocculation devices, and effects of particle size and ash content on the filtration rate. Detailed tests were performed on existing installations utilizing coal refuse filters to determine optimum flocculant dosage, solids concentration, and disc rotation speed to form a dischargeable cake. Information was also obtained showing effect of flocculant addition on the moisture of the filter cake. Our test work has shown that the continuous filtration dry cake rates for fine coal refuse correlates very well with the data developed and presented by Silverblatt and Dahlstrom1 except that the curves have moved up the scale by use of polymer flocculant additions. These polymers were not available at that writing. Finally, our tests also confirmed the general premise stated by S. K. Mishra2 that controlled additions of polymer flocculants will reduce filter refuse cake moistures somewhat.

Jan 1, 1974

By Toyohisa Fujita, Mitsuo Mamiya, W. Tai Yen, Hiroki Kogita

Green algae taken from an acidic mine drainage and blue-green algae taken from an alkaline hot spring stream were collected and tested for their ability to recover or remove dilute metal ions. Experimental results demonstrated that unwashed blue-green algae and washed green algae effectively adsorbed base metals ions and eluted them at pH 1. It was also found that washed and dried algae adsorbed precious metal ions more effectively than unwashed algae. For example, the washed and dried blue-green algae was capable of adsorbing 0.31 kg of gold per kg of algae. The gold from tetrachloroaurate solution which was adsorbed on washed blue- green algae was found to change to a metallic state following initial metal binding. In the case of a dilute gold complex solution leached with thiourea, only a small amount of gold could be captured by algae. Further experiments were conducted on components of the algae, such as alginic acid, agar, cellulose and chitin and mixtures of these components, in order to determine their contribution to metal adsorption characteristics. However, a mixture of these two components demonstrated both good adsorption and desorption characteristics indicating an interaction between the individual components.

Jan 1, 1995

By K. A. Natarajan, Abha Kumari

In this paper, electroleaching and electrobioleaching of ocean manganese nodules are discussed along with the role of galvanic interactions in bioleaching. Polarization studies using a manganese nodule slurry electrode system indicated that the maximum dissolution of iron and manganese due to electrochemical reduction occurred at negative DC potentials of -600 and -1,400 mV(SCE). Electroleaching and electrobioleaching of ocean manganese nodules in the presence of Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans at the above negative applied DC potentials resulted insignificant dissolution of copper, nickel and cobalt in IMH2S04 and in sulfuric acid solution at pH 0.5 and 2.0. Mechanisms involved in electrobioleaching of ocean manganese nodules are discussed. Galvanic leaching of ocean manganese nodules in the presence of externally added pyrite and pyrolusite for enhancement of dissolution was also studied. Various electrochemical and biochemical parameters were optimized, and the electroleaching and galvanic processes thus developed are shown to yield almost complete dissolution of all metal values. This electrobioleaching process developed in the laboratory may be cost effective, energy efficient and environmentally friendly. Keywords: Ocean manganese nodules, Electroleaching, Electrobioleaching, Galvanic interactions

Jan 1, 2003

By Martin G. Taube, Dwayne A. Lewis

Drainage improvements at O’Hare International Airport, Chicago, IL included the installation of a drainage and storm water system to control the over?ow from nearby Willow Higgins Creek. This involved construction of a weir structure at the creek and the channeling of water from the weir to a newly constructed reservoir via three, 3.7-m(12-ft) diameter, underground storm sewer lines. The sewer lines passed beneath an existing, 2.3-m (90-inch) diameter high-pressure water main, which was to remain in service throughout sewer line installation. The soil composition around the water main ranged from medium stiff clays to silty sands and sandy silts. Triple-fluid jet grouting was used to stabilize the variable soil profile beneath the water main in preparation for tunneling and installation of the sewer lines. The varying soil strata presented a challenge to the project team to establish a constant set of jet grouting parameters throughout the stabilized zone while keeping within the specified unconfined compressive strength range for the stabilized zone of 690 to 1,380 kPa (100 to 200 psi)at 28 days.

Jan 1, 2003

By C. G. Janes, N. Finkelstein, L. M. Jr. Johnson

In May, 1966, the General Services Administration (GSA) announced a program to encourage additional domestic production of copper in the interest of national security. Upon learning of the copper production expansion program, Duval Corporation entered into negotiations with GSA for development of the Sierrita copper-molybdenum property under the program. Agreements between the GSA and Duval Sierrita Corporation, a wholly owned subsidiary of Duval Corporation, were concluded in 1967, with the Sierrita Concentrator beginning operation in 1970 at a designed capacity of 72,000 STPD. Subsequent additions and modifications have expanded capacity to 90,000 STPD. This property, located approximately 20 miles south of Tucson, Arizona, represents an investment including working capital of about $200 million. These funds were provided from an $83 million loan from GSA, repayable in copper valued at 38 cents per pound, $48.75 million of bank loans subject to U. S. Government V- loan guarantees, and Company capital contributions. Current general operating statistics for the concentrator are listed below: Ore Grade 0.32% Cu., 0.030% Mo. Daily Tons Milled 90,000 STPD Annual Copper Production 94,600 Tons Annual Molybdenum Production 15,000,000 Lbs. Annual Silver Production 1,135,000 Ounces

Jan 1, 1976

By Robert A. Jankowski, John J. McClelland, Charles Babbitt

The purpose of this study was to develop guidelines to enable mine operators to maximize longwall dust reductions, while optimizing the use of ventilating air and shearer water consumption. A recent investigation by Foster-Miller Incorporated, under a Bureau of Mines contract, quantified the effect of increased face airflow and shearer water supply on face dust concentrations. A 70 pct reduction in respirable dust levels was achieved at the shearer operator's position during tail-to-head cutting by an incremental increase in average face air velocity from 83.8 m/min (275 fpm) to 123.4 m/min (405 fpm), while holding shearer water consumption constant. A 60 pct reduction was achieved through an increase in shearer water pressure and flow from 9.1 kg/cm2 (130 psi) at 0.19 m3/min (50 gpm) to 19.7 kg/cm2 (280 psi) at 0.28 m3/min (75 gpm), holding airflow constant. The combined effects of increasing both parameters produced an 85 pct reduction. Optimal usage of ventilating air and shearer water consumption can be achieved by defining a "point of diminishing returns" above which further increases no longer reduce dust concentrations, proportionately .

Jan 1, 1987

By M. G. Bakker, D. D. Murphy, D. R. Spears

The US Bureau of Mines studied the mechanism of shear flocculation of silica with stearic acid. At pH 6 and 12, shear flocculation of silica was not observed with the addition of up to 4.5 x 10-4 M (5 lb/st) stearic acid. However, in the presence of 0.1 M NaCl at pH 6, shear-flocculation was observed at stearic acid concentrations = 10-4 M (22.5 lb/st). Adsorption of stearic acid on silica was studied using stearic acids spin-labeled with a doxyl group. The presence of the spin-labeling group did not greatly affect the shear flocculation properties of stearic acid. Adsorption of labeled stearic acid on silica was measured as a function of surfactant concentration with electron paramagnetic resonance (EPR) spectroscopy and gas chromatography (GC). The GC results indicated that the stearic acid was predominately adsorbed on the surface at concentrations above 10-4 M. From the EPR spectra, the formation of aggregates of stearic acid was apparent at concentrations of approximately 10-5 M (22.5 lb/st). The results are discussed in terms of a possible mechanism for shear flocculation.

Jan 1, 1995

By F. Chavan

1. Introduction The new Cleuson-Dixence development project in the Valais region of the Swiss Alps, involves the construction of a new link between the Grande-Dixence dam (still world recordholder with a height of 300 m) and a power house on the banks of the Rhone consisting of approximately 20 km tunnels and shafts, and involving a fall of about 1900 m. This 1200 MW project will replace an existing 900 MW structure built in the 1960s, no longer considered efficient. The new project comprises a head race drift about 15.8 km long, a surge tank, an inclined cascade shaft 4,2 km long, and an underground power station. The work for lots C and D (head race drift from Tortin to Tracouet, surge tank, six adits ans the complete inclined shaft) was awarded to our group (Prader, Locher, SELL, Losinger, Evéquoz, Dénériaz, Imboden and Bouygues). The geological conditions are: Lower part of the shaft (300 m) : sedimentary - sandstone and soft tertiary sandstone. Middle and upper part (2600 m) : metamorphic - upper and lower carboniferous, interrupted by triasic chipping. The horizontal section (7 km) consists of mica

Jan 1, 1995

By James T. Furey, Laurence D. Hartzog, Graham A. Karklin

A pilot plant study was conducted to evaluate the performance of column flotation in the recovery of gold from the gold/tellurides bearing sulphide ore of the Kensington deposit (Echo Bay Mines and Coeur d'Alene Mines), and to compare the performance of column flotation to conventional "mechanically agitated" flotation, Tests were conducted in a 150 mm diameter, 60 liter glass column with ground flotation feed split from a concurrently operating conventional pilot plant, thus permitting direct comparison between the two processes. During the test program, the effects of wash water rate, gas holdup, froth depth, residence time and flow bias were examined in detail. The bulk metallurgical ore sample had an average "as tested" feed grade of about 5 git gold and 1 % sulphur, with the gold occurring primarily as inclusions of montbrayite (Au2Te3) in pyrite, with minor amounts of native gold. Conventional flotation using 60 minutes residence time across four roughing stages recovered about 97% of the gold into concentrates grading about 50 g/t Au. A single stage of column flotation roughing using 25 minutes residence time matched the conventional circuit recovery of97%, but produced grades of +150 git Au.

Jan 1, 1992